The present invention relates to a method for producing concentrates from copper-bearing raw materials, such as ores.
For treating primary copper raw materials, there are mainly two principal lines. One is the concentration—smelting—electrolytic refining line, and the other is leaching, such as the heap leaching—liquid-liquid extraction and electrolytic recovery line. With respect to reasons connected to raw material quality, environmental protection, geography and economy, both processing lines are meeting growing difficulties.
When starting to concentrate copper-based raw materials, we often face a situation where a large share of the mineralization is oxidized and possibly difficult to flotate. Among these are particularly copper ore deposits containing copper silicates and iron oxides. Also mixed grains with copper sulfide and pyrite may be nearly impossible with respect to flotation. A specific group of problems is represented by finely divided, often pyritic copper-zinc-lead ore deposits. The treatment of said ore deposits by traditional methods usually renders a fairly weak result as regards yields and concentrate contents. When transport costs to the smelter often are too high with respect to competition, even with a high-quality concentrate, they are even more so with a low-quality concentrate. What is more, in that case environmental hazards are increased at two separate locations, for instance because of arsenic. The smelting process itself typically includes many steps, among them smelting for example in a flash smelting furnace, converting, anode furnace treatment; sulfuric acid production for gases, and electric furnace or concentration process for slag. Often the reason for multistep smelting processes that are economically ineffective is the poor quality of the feed, i.e. the concentrate.
As regards the second prevailing method—processing based on heap leaching—it is likewise facing harder times. As long as the ore neither contains remarkable amounts of precious metals nor remarkable amounts of copper as chalcopyrite, CuFeS2, or as some other compound that is hard to dissolve, the situation is fairly good. However, as a rule, a growing share of raw materials even in already functioning mines is particularly formed of slow-dissolving copper minerals. This means increasing expenses. Another drawback of the method based on liquid-liquid extraction is the restricted lifetime of nearly all mines. If the whole process chain from the mine to cathode copper is based on one deposit only, the plant generally faces an unsound situation, as the volume of the ore body is gradually used up. As a result, the rate of profit for the invested capital is not optimal.
In geology, it was found out already at least a hundred years ago that metal sulfides tend to turn, for instance when precipitating from a solution ions of another element to sulfides. The observations gradually accumulated into real knowledge of the reasons of this phenomenon, to the extent that roughly 50 years ago, a patent U.S. Pat. No. 2,568,963 was published on the matter. According to said US patent, copper concentrate is divided into a fraction to be leached, and into a fraction used in the precipitation of copper sulfide (CuS). The obtained CuS is leached into sulfate in order to produce copper. The solid and soluble side components are simply removed from the process. Later, in 1956, the same inventors published a new patent, U.S. Pat. No. 2,755,172, where the metal ions of the solution, i.e. copper, cobalt, nickel and zinc, are precipitated in succession as sulfides, in the order CuS, CoS, NiS, ZnS, by using a metal sulfide of the MeS type that is more soluble than the element to be precipitated. In the precipitation process, the pH gradually rises, so that for instance in the precipitation of zinc sulfide (ZnS), the pH of the sulfate solution is within the range 6.2-7.
Because the starting point in the method of the U.S. Pat. No. 2,755,172 is the leaching of the raw material resulting from the production of sulfuric acid, the employed pH range 6.2-7 means that there is an economically demanding neutralization step. This fact is emphasized even further, when the suggested neutralization reagents are among others ammonia, lye or Ca(OH)2, or when a suggested sub-step of the process is a reaction where Fe3+ is reduced by hydrogen sulfide (H2S), producing sulfur, Fe2+ and H2SO4.
The weakness in the know-how of the processes described in the above mentioned U.S. Pat. Nos. 2,568,963 and 2,755,172, as well as the both chemically and economically unrealistic approach, are now, almost 50 years later, revealed by several features of the above mentioned US patents. First of all, in reality the natural sulfide minerals are not mainly of the type MeS only, but their metal/sulfur ratio (Me/S ratio) fluctuates within a wide range. Several metal sulfides are alloyed sulfides in the significance that metal (Me) is partly replaced by other sulfides, for example sulfur is replaced by arsenic and antimony, not to mention precipitation grains and other structural impurities, in comparison with pure MeS-type model minerals. As a consequence of the above mentioned facts, the method according to the U.S. Pat. No. 2,755,172 simply does not work with real raw materials. The method according to the U.S. Pat. No. 2,568,963 has better chances to function, but it does not offer a solution for example how to handle iron balances and acid balances. In addition, the U.S. Pat. No. 2,568,963 states that copper concentrate is needed in the leaching process, because other concentrates are too poor for leaching. What is more, a commercial-quality metal copper product cannot be achieved by the method according to the U.S. Pat. No. 2,568,963.
One reaction type in the production of rich copper concentrates is:CuFeS2+Cu2+═CuxS+Fe2+  (1)
The reaction (1) has often been found as slow. Therefore a solution has been searched in the direction of reducing conditions. The employed reductants have been for example elemental copper (Cuo), chromium (Cro), zinc (Zno), cobalt (Coo), nickel (Nio) or iron (Feo), sulfur oxide (SO2) or organic reductants. In laboratory conditions, the obtained reaction time for the reactionCuFeS2+Meo+Cu2+→CuxS+Fe2++  (2)is one hour, but it is understandable that in reality, a contact for example between Feo powder and CuFeS2 grain is not easily maintained. As such, the main principle itself for using metal powder is, for economical reasons, impossible in commercial processes. As for the use of SO2, it results in an excess of H2SO4 acid that is created in the process.
Moreover, it has been found that for producing copper-rich copper sulfide (CuxS), there are in principle two ways, i.e. a conversion based method according to reaction (1), and a selective leaching route by using an acidic reagent. The latter can be illustrated for instance by the reaction (3):1,8CuFeS2+4,8O2+0,8H2O═Cu1,8S+1,8FeSO4+0,8H2SO4  (3).
Thus a typical process based on selective leaching produces a remarkable amount of sulfuric acid and problematic FeSO4 solution, without essentially increasing the usage value of the copper sulfide product, because it contains harmful ingredients, such as FeS2 and silicates.
According to the DE patent application 2,207,382, CuFeS2 concentrate is treated in the presence of copper sulfate (CuSO4) by conversion in the temperature range 90-180° C. into CuxS and FeSO4. The obtained FeSO4 solution is hydrolyzed in an autoclave in the temperature range 180-230° C. into a solid Fe3+ compound and H2SO4 solution. The solid copper sulfide (CuxS) is leached by oxidizing with H2SO4 into CuSO4, which after cementing purification carried out by elemental copper (Cuo) is reduced into copper with hydrogen. The method according to the DE patent application 2,207,382 is feasible with pure concentrates that contain only small amounts of for example zinc, lead and pyrite. Similar problems are also included in methods described in the patents CA 1,069,317 and U.S. Pat. No. 3,957,602. In the former method, CuFeS2 concentrate is converted to CuxS and FeCl2 solution by chloride leaching. CuxS is leached, and after cleaning, metallic copper is reduced from CuCl. Impurities are removed from the FeCl2 solution, and by means of electrolysis, FeCl2 is turned into FeCl3 solution and metallic iron. This method could be fairly feasible, if neither the purity of the product nor the economical values in particular would have any importance. The method according to the U.S. Pat. No. 3,957,602 is a basic version of two main lines based on the production of CuxS by using fairly pure copper concentrate. Here the iron contained in CuFeS2 is in connection with the leaching of CuxS turned into jarosite. However, the method according to the U.S. Pat. No. 3,957,602 does not take into account for example the recovery of precious metals and MoS2, but its use brings along additional expenses in comparison with existing mainstream methods.
Nearer to the method of the present invention come the processes described in the reference publications Yuill W. A. et al, Copper Concentrate Enrichment Process, SME-AIME Annual Meeting, Los Angeles, Calif., 26 Feb.-1 Mar. 1984 and Bartlett R. W., A Process for Enriching Chalcopyrite Concentrates, New Orleans, 2-6 Mar. 1986, pp. 227-246. As for the first alternative, written in 1984 by Yuill et al, the most serious drawbacks are connected to the leaching process carried out at the temperature of 200° C. and to the oxidation of nearly all sulfidic sulfur, and to a great extent also of pyritic sulfur, into sulfate, i.e. into sulfuric acid. The situation is attempted to be improved by the use of lime both in the leaching autoclave and in the conversion autoclave, and also in the copper removal of the solutions. The obtained conversion product is further subjected to flotation, which causes extra expenses, and in reality also copper and precious metal losses in this process. As a whole, for the obtained rich CuxS product, there is not found a further treatment process that would be more advantageous with respect to usage, but the CuxS concentrate must compete with the traditional copper concentrate.
In the process described by Bartlett in 1987, the autoclave steps are combined into one CuFeS2 leaching conversion step operated at the temperature of 200° C. From the point of view of the equipment technology, the process is simplified. Still the problems related to the creation of H2SO4, to iron removal and the use of lime for the most part remain the same. As a back balance for the simplification, the degree of conversion of the CuFeS2 concentrate is essentially weakened, mixed FeS2-copper sulfide grains remain in the product, and the recovery of copper in the final concentrate is lowered for instance owing to increased problems in the selective flotation of the end product. From the point of view of the smelter, the obtained product is still not attractive in comparison with the traditional concentrate.